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One of the major unknowns in the use of iron and copper is the quantity of various low level impurities, particularly hydrogen, that may be introduced during the deoxidizing process prior to casting. Introduction of hydrogen results in the embrittlement of components made from these metals. In addition, hydrogen bubbles in castings of copper result in the formation of blisters and swelling of the casting. During the last year, a significant effort has been made at the University of Missouri-Rolla to understand the permeation of hydrogen in galvanized automotive steel components and the hydrogen absorption during the electrodeposition of copper. The galvanized steel components consisted of automotive upholstery clips, seat belt holders and brake clips. Although the process of galvanization protects the steel from corrosion the process invariably causes the steel to pick up enough hydrogen to make it susceptible to brittle fracture.

Jan 1, 1992

By W N. Qin, B Zhang, W Z. Huang, X S Mo

Mengzi Mining and Metallurgy Co, Ltd (MMMC) polymetallic ore contains recoverable amounts of tin, lead, zinc, pyrite and other valuable minerals. With the change in orebody, the run-of-mine (ROM) ore changed from lead zinc rich ore to tin rich ore, where the current ROM ore contains 0.45–1.25 per cent tin, 20–25 per cent sulfur, and 0.20–0.35 per cent copper. The previous processing flow sheet was designed to recover lead zinc rich ore and it is clearly not suitable for recovery of the current tin rich ore, where the tin was already overground and tin recovery was only 50.19 per cent. The Mengzi Mining Company has recently modified the concentrator circuit by incorporating rod mill and high efficient Derrick fine screen. As a result of the improved classification efficiency, the overgrinding of the tin ore was minimised and the recovery of tin was improved by 20.71 per cent, ie from 50.19–70.90 per cent. An additional $10 M revenue per annum can be achieved as a result of the modification.CITATION:Qin, W N, Huang, W Z, Mo, X S, Zhang, B and Wheeler, J, 2014. Application of high efficient fine screen to improve the recovery of tin concentrate, in Proceedings 12th AusIMM Mill Operators’ Conference 2014 , pp 487–492 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Sep 1, 2014

By Sergio Romo, Gerardo Fuentes, Adriana Garcés

Refractory ores with gold entrapped in a sulfide matrix, must be processed in a first stage with chemical or physical treatment in order to liberate the gold for a subsequent step with cyanide. The goal of this research was to compare the refractory behavior of a gold concentrate treated using two pretreatment methods: pressure oxidation and fine grinding. Gold concentrate from Viacha Mining in Bolivia, assaying 50.11 gpt, with a P80 of 131 ?m was used. The concentrate contains arsenopyrite and pyrite as main sulphides in the matrix. A preliminary test with cyanide resulted only 13.5% of gold extraction with 4 hours of stirred leaching. Using an autoclave oxidation with sulphuric acid and 12 bar oxygen pressure at 220°C pretreatment, the cyanide gold extraction increased to 69.0% in 4 hours. When material was ground in a pulverizer, producing a P80 of 20 ?m and the product directly cyanided, stirred leaching gave a gold extraction of 65% in 4 hours. For both techniques, the effect of temperature, sulfur produced in autoclave oxidation, optimization of the cyanide and the nature of residues were also studied.

Jan 1, 2014

By John E. Urosek, Dennis A. Beiter, William J. Francart

In This Chapter [Designing bleeder systems Examining and maintaining bleeder systems Evaluating bleeder system effectiveness] INTRODUCTION Bleeder systems are that part of the mine ventilation network used to ventilate pillared areas in underground coal mines. Pillared areas are those in which pillars have been wholly or partially removed, including the areas where coal has been extracted by longwall mining. Bleeder systems protect miners from the hazards associated with methane and other gases, dusts and fumes, and oxygen deficiency that may occur in these mined-out areas. Effective bleeder systems control the air passing through the area and continuously dilute and move any methane-air mixtures and other gases, dusts, and fumes from the worked-out area away from active workings and into a return air course or to the surface of the mine. A bleeder system includes the pillared area (including the internal airflow paths), bleeder entries, bleeder connections, and all associated ventilation control devices that control the air passing through the pillared area. Bleeder entries are special air courses designed and maintained as part of the mine ventilation system. [This chapter focuses on the design, examination, maintenance, and evaluation of bleeder systems in underground coal mines.] The history of coal mine explosions in the United States is a reminder of the importance of adequate ventilation. Some of those disasters were the result of inadequately ventilated pillared areas. The importance of developing bleeder systems to ventilate these pillared areas and evaluating the bleeder system’s effectiveness is reflected in present-day federal regulations. For more information on bleeder systems, see Tisdale [1996] and Urosek and Francart [2002].

Jan 6, 2006

By Charles E. Whieldon

Because of the importance of recovering more oil from known petroleum reserves, increased emphasis has been placed on the application of secondary-recovery methods. It is hoped that these reports will stimulate interest in improved waterflooding techniques, thereby discouraging premature abandonment of oil leases. Previously published reports3 4 5/ have described the results of water-flooding in Roane, Pleasants, and Wetzel Counties, W. Va. (fig. 1). This report describes the South Penn Oil Co. Yawkey waterflood property, Berea sand, Griffithsville oilfield, Duval District, Lincoln County, W. Va. (fig. 2).

Jan 1, 1961

By R. O. Suzuki, S. Natsui, R. Nashimoto, T. Kumagai, T. Kikuchi, H. Takai

A newly developed computational fluid dynamics model based on a multi-phase particle simulation is presented for predicting the entrainment behavior of liquid metal into slag due to rising single bubble. By comparing results calculated using this model against experimental data, it was found that the transient behavior of bubbles and the two immiscible liquids can be accurately estimated by this method. The rupturing of the thin water film surrounding the bubble was less reliably predicted, but this 3-dimensional unsteady numerical model still nevertheless provides valuable new information for directly predicting the change in the liquid-liquid interface over time. This model predicted the Young-Laplace pressure, and the interface shape was simulated with volume conservation using this model as opposed to the conventional particle method. This simulation model can calculate both of dispersed phase and continuous phase seamlessly.

Jan 1, 2015

By R. Horn

The financial performance of the mining industry, measured by return on equity and capital employed, has been inadequate for many years. Consequently, mining companies are not valued highly by stock markets. Among the more important causes for the decline in the sector are the conservative working practices and reluctance to adapt to change or to develop new technology, although some relatively recent developments such as SX-EW copper and heap leach gold have made a significant impact on the cost base of these metals. The subsidiary business of metals exploration suffers from the same problems as the rest of the industry. The rate of discovery of ore deposits has been falling over many years and since the late 1990s, there has been a sharp reduction in global exploration activity that, if it continues, will further reduce the global supply of new ore reserves. New technological developments, particularly in geophysics and different operational practices may help reverse this trend. Statistical methods for evaluating the risk of ore reserve estimates will also improve the economic return on new mining ventures. The science of ore geology has potential to reduce exploration risk, but modern ore genetic theory has not contributed greatly to recent discoveries. The junior exploration market has almost ceased, but a new style of technically specialized efficient small exploration company has emerged which, in alliances with major companies, may be the vehicle for future global exploration. Exploration in virgin terrains has a higher risk discounted potential for the discovery of very large orebodies than that in known mining districts but mining companies have little financial incentive for this high risk exploration because of the very long lead time to production. Alliances of two or more companies reduce the risk in both the exploration for and development of major new mining projects. If new orebodies are not found, then metal prices will inevitably rise in the long term. Governments of industrial countries, particularly those with a significant mining industry such as Canada, should therefore provide assistance to exploration in order to maintain their existing industry and prevent a metals shortage from developing in the future with its consequent impact on the world economy and increased environmental damage as large lowgrade orebodies are developed to supply an expanding market.

Jan 1, 2002

By Lucien Eaton

THE accuracy with which detachable bits can be sharpened and gauged by grinding has made possible refinements in design that cannot be attained by forging. This applies particularly to gauging, where accuracy is of the greatest importance. Gauge changes of 1/32 in., and even 1/64 in., are now feasible, whereas 1/16 -in. changes are about the limit in accuracy for forged bits. These small changes in gauge are feasible, however, only with certain fundamental principles of design. Within reasonable limits the depth to which a hole can be drilled is dependent primarily upon the loss in gauge, or diameter, of the bit. In drilling there is always a certain amount of "overdrilling"-the bouncing of the bit in the hole and its rotation wear away the walls-so that the actual diameter of the hole is greater than the diameter of the bit. The amount of this overdrilling depends upon the kind of rock being drilled, and is sometimes sufficient to permit the use of the same diameter of bit for the full depth of the hole. If, for instance, a 1 3/4-in. diameter bit overdrills 1/32 in. in diameter, and loses 1/64 in. in gauge in drilling 2 ft. of hole, the bottom of the hole is still 1/64 in, larger than the original diameter of the bit, and a new 1 ¾ -in. bit can be used to drill the next 2 ft. If, however, the overdrill is less than the loss in gauge, the next bit must be of smaller diameter. The first requisite in reducing loss of gauge in drilling is to have the points and corners of the wings of the bit cut circles of the same diameter, thus producing a reaming surface with continuous bearing on the wall of the hole. The greater the reaming surface, the less the gauge loss that is to be expected. If the reaming surface, however, is too large, it may impede the rotation of the bit through excessive friction. This principle was probably first recognized in the design of the Carr bit (Fig. 1), a bit with a single cutting edge of wide angle, a definite reaming edge, and a large center hole. The same idea was followed in designing the double-taper cross bits, which are now standard almost everywhere.

Jan 1, 1939

By David A. Newman

In the 1950s, an underground limestone mine in the Camp Nelson Formation was developed from an outcrop exposure. The Camp Nelson Limestone met the requirements for DOT stone and aggregate. The mine ownership changed several times, which resulted in distinct mine designs in Level 1. Based upon the production requirements and an attempt to maximize limestone recovery from a single level, a variety of pillar dimensions and three mining heights (28 feet, 68 feet and 83 feet) were developed. This was sufficient prior to reaching the mineable extent of Level 1. Level 2 introduced previously unknown ground control challenges. These included: ??Pillar spalling and a significant back fall on the Level 1 - Level 2 decline; ??Laminated immediate back and the need to bolt the back on Level 2; ??Delamination of the back adjacent to the Level 1 - Level 2 decline resulting in bolt shearing, and floor heave in isolated areas of Level 2; ??Multiple level interaction, the need to design pillars based upon numerical modeling as columnization between Levels is infeasible: and ??Water inflow as the mine approached river. To meet the challenges, a testing program to determine rock strength and physical properties was implemented along with numerical (LAMODEL and FEM) modeling, and borescope inspections of the immediate back. The efforts to implement rock mechanics were re-focused as Level 2 approached its mineable limit and the initial attempt to drive a decline to Level 3 was unsuccessful. A second decline site was selected based on inter burden stability, a consistent shale parting for the Level 3 back horizon, results from numerical models of the immediate back and multiple level interactions, and the projected low water inflow. The mine is operating on Levels 2 and 3 with long-term planning and ground monitoring for Level 3.

Jan 1, 2014

By Ray Holmes

This paper describes one approach in our search for greater stope efficiency and lower costs. The uppers technique with wagon drills in cut-and-fill stopes was commenced, on an experimental basis, in our Sudbury mines in 1961. To date, the method has developed into an accept-ed production operation. This technique minimizes the intermittent aspects of the drilling, blasting, mucking and filling operations associated with other cut-and-fill methods. However, studies show that not all cut-and-fill stopes are suitable for uppers, the governing factors being ore width, dip of the orebody, and ground conditions where large areas of unsupported back remain stable for ex-tended periods of time until rockbolting can be accomplished. Introduction F ALCONBRIDGE Nickel Mines Limited operates seven mines within a radius of 30 miles of the city of Sudbury, Ontario. These mines extend in depth from 1,500 to 6,300 ft. The ore consists of brecciated, massive and disseminated sulphides, and the ore minerals consist primarily of pentlandite, pyrrhotite and chalcopyrite. Widths are generally irregular and vary up to 150 ft. The ore dips at between 35 and 90 degrees.

Jan 1, 1965

By John Watson, Jay Rodgers

The world of blasting and explosive use continues to evolve into one which relies heavily on measurement, quantification, process control, and information feedback as a means of contributing to operational improvements. More than ever good blast performance, once thought to be a result of someone’s “artistic ability”, is giving way to “hard science” and measurement techniques. Variability in product performance and process control is becoming less tolerable with increasing environmental issues, regulatory pressure and public awareness.

Jan 1, 1998

By Marty Hudyma

There are five independent seismic source parameters: seismic energy, seismic moment, source size, location, and time. These parameters are investigated primarily for evaluating seismic hazard and understanding seismic source mechanism(s). Broadly, seismic hazard is the likelihood of occurrence of an event of a given magnitude. Seismic source mechanism is the mode of rockmass failure that causes the seismic event. Most mine seismicity analysis techniques use a combination of a few of the independent source parameters. In literature, time is perhaps the least discussed and most underutilized of these five parameters. This paper discusses three applications of time to understand both seismic hazard and seismic source mechanism in mines. The first application is time history analyses, which focuses on how other parameters of interest vary over time. Examples are presented to show how time history analyses can give information about both seismic hazard and seismic source mechanism. The second application is analysis of time-between-events. This technique can give insight into seismic source mechanism. The final application is phasor analysis using the Schuster test. This method can be used to investigate how seismic source mechanisms may change over time. For both time-between-events analysis and phasor analysis, there are important implications with regards to scale dependence and the breakdown of self-similarity of mining-induced seismicity.

Jun 1, 2012

By B. R. Mohal, F. F. Aplan, S. Chander

"ABSTRACTThe wetting behavior of coal has been determined in the presence of nonyl and octyl series of nonionic surfactants containing polyethoxy groups of different sizes. The wetting behavior was determined by a modified Walker technique and contact angle measurements. The results show that the wetting rate of coal varies with coal rank, structure of the hydrocarbon chain and number of ethoxy groups in the hydrophilic group of the surfactant. A close relationship has been observed between the wetting rates and the short-time advancing contact angles.INTRODUCTIONNonionic surfactants are used in many applications to improve wettability of solids. They can be used as dispersing agents [ 1 ] or dust suppressants [ 2-5 ] and to control the flotation, flocculation, filtration and flowability of particulates. This study was undertaken to study the effect of a series of nonionic surfactants on the wetting behavior of coal. Two series of surfactants, namely, ethoxylated nonyl and octyl phenols were used. The nonyl phenol series contains a linear hydrocarbon chain containing nine carbon atoms ( CH3( CH2 )8-C6H4,- ( OC2H, ) n, where n is the number of ethoxy groups) . The octyl phenol series contains a branched hydrocarbon chain containing eight carbon atoms ( ( CH3 ) 3CCH2C ( CH3)2-C6H4- ( OC2H4) n ).Initial wetting rates — modified Walker techniqueWalker et al. [ 6 ] developed a sink test to determine the wettability of coal particles. In this technique, coal particles are dropped individually onto the surface of surfactant solutions of different concentrations, and the most dilute concentration in which coal would sink ""instantaneously"" is determined. This technique is a variation of a procedure initially developed by Draves and Clarkson [ 7 to determine the effect of surfactants on wettability of cotton fibers. In this procedure commonly known as the Draves test, the concentration of a surfactant required to give a 'sink time' of 25 s is determined.Other investigators have used slight variations of the Walker method to study the wetting of fine particles [ 8-12 ] , by measuring the time for wetting a given amount of powder. There are differences in the manner in which the time of wetting is measured, however. Some investigators measure the time for the last trace of coal to disappear [ 8,11 ] whereas others determine the time for the vast majority of particles to sink [ 10 ] . Some investigators report a wetting rate [ 11 ] which is a value calculated on the basis of the time for the 'last traces of particles to sink'. The differences in choice of criterion for wetting can result in different values for wetting rates as illustrated in Fig. 1 which is based upon our measurements of the amount of powder wetted as a function of time. The procedure for wetting rate measurements is described in the next section. The technique in which time for disappearance of the 'last traces of particles' is determined gives the smallest value for the wetting rate, while the technique in which the time for wetting of the 'vast majority' of particles is determined will give intermediate rates. The terms 'last traces of particles' and 'vast majority of particles' are usually not defined explicitly and can lead to large differences in wetting rates measured by different investigators. Since our measurements show that the wetting rates for an assembly of powder decreases with time, we have used the initial wetting rates to compare the wetting behavior of coal in solutions of different surfactants."

Mar 1, 1989

By John E. Riddell

This review is based on a questionnaire sent out to 24 Canadian exploration companies, known to be active in the mineral exploration field, concerning their practice in the use of applied geochemical techniques in Precambrian terrains. Of the companies questioned, 16 returned completed questionnaires and of these, ten companies stated that they had some experience in this type of work. Of the companies who did not answer or return the questionnaire in time I know that at least three have been quite heavily engaged in applied geochemical work in the glaciated Precambrian areas. I also know that these three companies have been reasonably successful in their ventures, so I feel the following figures and opinions are, if anything, on the pessimistic side. Before continuing with the paper I would like to express my personal thanks to the Chief Geologists, Geologists and Exploration Managers of the various companies who cooperated in making this review possible by answering the questionnaires and making available summaries of their work in the glaciated terrains. Of the companies who stated that they had no experience in geochemical prospecting in glaciated Precambrian areas it is interesting to note that only two have reported that they do not use applied geochemistry at all. Two companies have reported that they make use of applied geochemistry in New Brunswick in areas of-light glaciation. Their opinion is that it is a useful aid to other techniques in evaluating mineralized areas.

Jan 1, 1960

The distribution or the userul minerals or Tasmania is so closely connected with the geological structure of the island that it will be convenient to briefly recapitulate the principal ract or its geological history, before describing the ore and other deposits. In many cases these present features which are very puzzling in the absence of knowledge of the structure of the districts in which they are found and in all cases it is of benefit to mining engineers to be acquainted with the main facts of the geological·history of the regions in which they have to carry on operations. Moreover, in speculating as to the origin of lodes and other deposits, we have constantly to rerer to the crumplings or the earth's crust by mountain-making movements, and to the igneous rocks which have been thrust throngh the other strata, to explain the origin or the lode fissures, and this discussion cannot be carried on without knowledge of local geology.The oldest rock yet met with in Tasmania are of Archooan age, and consist of highly crystalline quartzites and micaceous schist. They are most extensively developed towards the western side ot the island, forming a belt parallel to the West Coast from South-West Cape to Cradle fountain; but they also crop out here and there to the west and northward of this belt, where the covering strata or later age have been denuded away. Though not seen in situ in the eastern part of the colony, the abundance or pebbles derived from this formation in the conglomerates of the coal measures shows that it must in those day have formed land surfaces there also, and makes it likely that the archooan strata exist at no great depth: below many of the beds of later age now occupying the surface. Lying unconfortably upon these highly metamorphic strata we find next an extensive series of subcrystalline metamorphic conglomerates, grits, sandstones, slates, and limestones,of Cambro-Silurian age. These are well seen in the Gordou River Basin, along the West Coast Range, aloug the North Coast, and in the goldfields of the north-eastern portion of the colony. They are much broken by later intrusive rocks, and covered by younger rormations, but the series is spread more or less over the whole island.

By L. M. Lopez, J. A. Sanchidrian

This work investigates the performance of some of the measurement techniques used in vibrations from blasting monitoring. Two tri-axial geophones with cases of different size, shape and mass were tested in a shaker table. The geophone cases were mounted in three different ways on a granite slab fixed to the plate of the shaker: freely placed, sandbagged, and anchored. A known horizontal periodic motion of constant amplitude, and frequency from 16 to 200 Hz, was applied to the slab. Two amplitude levels of 5 and 20 mm/s (0.2 and 0.79 in/s) were applied when the sensors were sandbagged or anchored and only 5 mm/s when no holding force was applied. The transmissibility of vibrations from the rock to the longitudinal geophone as function of the frequency (i.e. ratio of response of the geophone to the slab motion) has been obtained for each measuring condition tested. The transmissibility is considered a composition of the transmissibility of the rock-to-sensor coupling and the transmissibility of the sensor’s mechanics and electronics. The transmissibility of the coupling is obtained normalizing the total transmissibility with the transmissibility of the seismographs, obtained by a suitable test to which a coupling transmissibility of 1 is assigned. Geophone-to-rock coupling is a resonant phenomenon that may alter the amplitude of the measured waveforms by a factor of 1.2 to less than 0.6. The most accurate and precise measurements are obtained with geophones anchored to rock. On the contrary, freely placed or sandbagged sensors are non-predictable planting conditions that lead to high amplification or damping in most of the frequency band analyzed. It is also shown that existing acceleration criteria to select coupling methods do not necessary lead to good quality (accurate and precise) vibration data.

Jan 1, 2014

By J. H. Tu

The technical paper correctly points out that the kriging variance is not a good measure of the uncertainty of the estimated (i.e., kriged) value of individual blocks. The authors claim that their proposed jack-knife method, which is a rekriging of each block by eliminating, in turn, one sample from m the original sample set and then taking the average of the rekriged estimates, not only gives good block estimates, but the resulting jackknife kriging standard deviation is a useful indicator of the "true uncertainty associated with block estimates." However, they immediately abandon the idea of using the block-by-block standard deviations, reasoning that these standard deviations are not independent and that there is no easy way to utilize them. There may be another reason for not using them. The jackknife standard deviations for individual blocks given in their example are mostly in the range of 0.004 to 0.005 oz/st (0.14 to 0.17 g/t) with only one block having a high value of 0.012 oz/st (0.41 g/ t). These individual block standard deviations are as low as the jackknife standard deviation for the mean grade of the entire shape, i.e., 0.0041 oz/st (0.14 g/t). Do they represent the "true uncertainty" .of the individual block estimates? Could the authors explain this? In a global shape consisting of a large number of blocks, any given sample will affect the kriged estimate of only those few blocks within its vicinity. This is the rationale for the authors' selective rekriging, making the jackknife algorithm more efficient. On second thought, why not do away with jackknifing altogether? Just cumulate and normalize, if necessary, the kriging weights of each sample used during the ordinary block kriging process, and then compute the global variance from these kriging weights and their respective sample grades? After all, isn't the global mean grade nothing but the weighted average of the samples used in the estimation? Reply by G.S. Adisoma and M.G. Hester The jackknife is one of the many tools in a practitioner's toolbox to solve estimation problems. The strengths of the technique lies in its simplicity, i.e., it uses the concept of mean and standard deviation and the fact that it can be easily combined with other tools, in this case kriging. Because the jackknife kriging (JK) estimate is also the mean of the pseudovalues, the JK standard deviation is attractive just as the standard deviation of the mean explains the variability of the data. The difference is that the pseudovalue calculation in jackknife kriging uses the ordinary kriging (OK) weighting scheme instead of simple arithmetic averaging. The data used to illustrate the jackknife technique in the paper con¬sist of high values that are roughly three times the low values. The resulting JK estimate of the block grades show that the highest estimate is roughly twice the grade of the lowest estimate. The contrast between the low and the high estimate is more evident in the JK estimate than in the OK estimate, even though the mean grades of the blocks for the two estimates are very similar. Nonetheless, in this paper, we are concentrating more on the need for a more realistic measure of uncertainty, or precision, for the estimate. Unlike its OK counterpart, the JK standard deviation of the blocks clearly reflects the original data variation. The highest JK standard deviation of the blocks is three times its lowest value. This follows our intuition that, when the samples used to estimate a block is more variable, the resulting estimation variance (or standard deviation) should be higher than the case where the samples are more uniformly valued. However, block-by-block standard deviation or variance is of little practical value in reserve estimation and classification, as well as in mine planning. One is usually more interested in quantifying not the variance of the individual block estimate, but the uncertainties associated with a much larger dimension, such as the minable reserve. Thus, the thrust of the paper is to find a simple way to obtain a single estimation variance or standard deviation associated with the reserve grade estimate. The discussion by J.H. Tu did not mention how one would obtain the global variance from the OK weights and the sample grades. As a technique that offers a data value-based measure of uncertainty for its estimate, the "leave-one-out" jackknife fills this need nicely through the block kriging shortcut approach described in the paper. Note: The first column and the last two columns of Table 3 in the paper should have contained a single number each, namely, an OK estimate of 0.0317, a JK estimate of 0.0333 and a JK standard deviation of 0.0041 oz/st, respectively, for the shape, as are obvious from the text.

Jan 1, 1997

By A. A. Toprac

The recurring problem of joint failures in ofjshore structures, coupled with the catastrophic collapse of many of these structures during Hurricane Hilda (1964), hay forced reconsideration of techniques presently used to design sfuctural joints. This paper presents results of studies conducted for this reconsiderat?on. Only right-angled T-joint,s and cross-joints where a brace was welded directly to the undisturbed surface of another tube (the chord) were considered. Tests were performed on seven T-joint and 12 oross-joint specimens of two types which were designed to give a broad but practicd range to the geometrical parameters believed to be of significance to the problem. These paramefers were expressed in dimensionless form. Each specimen of the T-joints was instrumented wirh both electrical resistance strain gauges and deflection dials. After the elastic tests were concluded, each specimen was loaded to failure by applying an axial tensile load to the brace. The cross-joints were tested to failure without measuring elastic strains. Theoretical solutions for the stresses in a cylindrical shell due to a distributed radial load over portions of its surface were numerically evaluated. The experimental results are compared with calculated values obtained using Bijlaard's and Dundrova's theoretical analyses. INTRODUCTION During the last IS years a large number of structures were built for the offshore exploration of oil both in the Gulf of Mexico and dong the Pacific Coast.' In all these offshore platforms, the tubular shape was extensively used since this shape provides outstanding strength in proportion to its weight, compared to other shapes, when all directions (x, y, z) are considered. Past use of tubes was hampered because of connection details. But with the advent of welding, it became no more difficult to connect tubular shapes than to connect rolled shapes. However, some problems have been encountered in actual construction in making welded joints that are adequate and yet economical. In trying to connect tubular members, designers have used a varied number of methods; consequently, some of the joints seem to be very expensive in detail while others appear weak and underdesigned. An inspection of offshore platforms owned by different companies indicates the use of a large number of joint arrangements and configurations. Such a variety of joint types in current service indicates the wide divergence of opinion that exists on how such joints should be designed, stiffened and strengthened. Fig. 1 shows the various types of connections that have been used in practice. This paper reviews the present state of art on tubular joints of the T and cross type, and presents results of experimental and analytical work being conducted at The U. of Texas and elsewhere. It is believed that such information will be of great help to designers for the formulation of proper design procedures for joints. CONNECTION PROBLEM Recently, cracks have been observed in several welded connections in offshore structures. Fortunately, until about 1 1/2 years ago, these failures have been localized and have not resulted in the collapse of structures or loss of life. For example, one platform installed in Louisiana during 1955 in about 60 ft of water had been subjected to four tropical hurricanes. This platform successfully withstood these forces and was later removed because operations at this location had been completed. Many joint failures were observed during the removal.

By William G. Wilson

In the U.S. today a number of tentative synthetic fuel plants have been sited by a number of major corporations. Each of these for a number of reasons have deferred construction. The only synthetic fuel plants under active development are those being built with the Energy Research and Development Administration (ERDA) under the authority of the Non-nuclear Energy Research and Development Act of 1974 (Public Law 93-577). Under this Act in joint sponsorship with industry, a series of coal conversion pilot and demonstration plants are being built to produce environmentally acceptable fuels from high sulfur coals. Before entering into a discussion of the factors associated with the siting of the ERDA coal conversion program, the difference between a pilot plant and a demonstration plant should be explained. The terms pilot plant and demonstration plant iridicate a size of plant rahter than technology. A pilot plant will vary from an "add-on" to an existing facility to an independent unit. Demonstration plants are a wholly independent, self contained plants, capable of independent operation to demonstrate the environmenta1, economic, and technica1 feasibility of a technology. These plants are from one-third to one-tenth the scale of a commercial operation and often described as plus 1000 tons per day of coal consumption. Their purpose is stated in the Act (Section 6, (3)(E),) "to accelerate the commercial demonstration of technologies for producing syncrude and liquid petroleum products from coal: Provided. That the Administrator shall invite and consider proposals from potential participants based upon Federal assistants and participation through guaranteed prices or purchase of the products .... " In site selection and plant design it has been suggested that special interest groups should be allowed to dictate the site and specify the design. Under the present program the industrial partner who is contributing 50 percent

Jan 1, 1977

The Khanong supergene-enriched copper resource, (~23 Mt at 3.0 per cent Cu), is one of several resources recently discovered in the intrusive-centred, sediment-hosted Au-Cu Sepon mineral field located in southern Lao PDR. The unusual style of hypogene mineralisation and unique set of conditions have resulted in the development and preservation of a supergene-enriched profile in this subtropical climate. Hypogene sulphide-rich mineralisation replaced calcareous cataclastite (fault rock) developed in the flat-lying, extensional, Khanong Fault. Spatially, the hypogene mineralisation was immediately out board of the skarn û hornfels Cu-Bi front and in board of sediment hosted Au-(As-Sb) mineralisation in a zoned intrusive-centred hydrothermal system. In situ supergene enrichment of the sulphide protore has resulted in a high-grade chalcocite-clay zone (~10.3 Mt at 4.6 per cent Cu) overlain by leached limonite-clay. A very high-grade in situ to exotic limonite-clay Cu-oxide-mineral zone is locally developed above impermeable dolomite at the base of chalcocite-clay zone. Exotic Cu-Mn-wad mineralisation developed for up to 1 km down the hydraulic gradient from the chalcocite-clay zone. Unique supergene enrichment conditions have resulted in the high-grade Khanong Cu resource in a subtropical weathering environment. The base of the currently active supergene profile coincides with the base of a flat-lying, pyrite-rich fault rock, dipping parallel to the hill slope, maximising Cu enrichment. Meteoric water has been channelled along subvertical faults and narrow horizontal subsurface zones. Low permeability of the fault host rock and footwall rocks has controlled the low volume and slow movement of groundwater required for supergene enrichment. Erosion has been slowed by the resistive ironstone and limonite-clay layer at surface, which assists in preserving the gentle slope over the chalcocite-clay zone.

Jan 1, 1999